Method of recovering mineral values



July 14, 1959 R. F. MocULLoUGH ET AL .2,894,809

' 'METHOD oF REcovERING MINERAL VALUES Filed July 6, 1955 2,894,809Patented' July' 14,1959` United States `Patent .t ice 2,894,809 METHODOF RECOVERING MINERAL VALUES Robert F. yMcCullough, Glenview, Ill., andJoseph B.

Adams, Lakeland, Fla., assignors, by mesne assignments, to the UnitedStates of America as represented by the United States Atomic EnergyCommission n Application July 6, 1955, Serial No. 520,376

13 Claims. (Cl. 23-14.S) n `Y This invention relates to the recovery ofmineral values from low grade phosphate-bearing ores. More partic11-larly, it relates to the recovery of phosphorus, aluminum and uraniumvalues from leached zone material from the Florida pebble phosphatefields. v

Processes for the recovery of P205 values from phosphate-bearing orescurrently in commercial use generally involve conversionV of phosphatesto water soluble forms by reaction of the ore with sulfuric acid eitherin a wet slurry system or a system involving the formation of anacid-mix which sets to relatively dry solid form and leaching the drysolids to recover the water soluble reaction products. The solution orextract is `generally treated with lime or other reactants to recoverphosphate salts or reacted with, for example, phosphate rock to formproducts such as triple superphosphate.

These processes all have two disadvantages in common. The processes allrequire a relatively large amount of acid in excess of thestoichiometric requirements for reaction to eiectually solubilze al1 themineralv components. In addition, the `slurry mixtures formed duringprocessing present ditlicult filtration problems.

In those processes Where recovery of minerals other than phosphorus fromthe low grade ore was attempted, the processes have shown the tendency,namely, that when conditions were set for high recovery of one componentonly, recovery of other components followed to about the same degree. Inother words if 90% uranium recovery was desired, suicient acid must beadded to solubilize about 90% of the phosphorus and other reactablevalues. f

It Iis a primary object of this invention to overcome the shortcomingsand disadvantages of the processes heretofore in use. v Y

It is an object of this invention to provide a method for recovery ofmineral values from low grade ores as represented by leached zonematerial, kaolinite ore and the like. Y

It is another object of this invention to provide a method forrelatively high recovery of all constituents convertible to watersoluble reaction products by reaction with sulfuric acid. l

It is a further object of this invention to provide a method forimproved solubilizing of uranium constituents of low grade phosphateores per unit of reagent used while depressing the reactivity of thephosphorus and aluminum values. Y

It is still another object of this invention to provide -a method forrecovery of constituents of low grade ore wherein the acid mix is heattreated thereby improving the filtering characteristics of the slurries.

These and other objects of the invention will be apparent `to thoseskilled in the art from the following description. I

' Briey, the instant invention comprises acidulating the low gradephosphate-bearing ore with sulfuric acid under conditions of strongagitation, heat treating the acid- Vmix at temperatures in the rangebetween about 200 C. and about 700 C., leaching the water soluble reac-70 2 1 .c tion products from the heat treated mix and processing theleached solution to recover at least one of the phosphorus, aluminum,uranium or other constituents.

More in detail, the following discussion is made relative to leachedzone material as a representative low grade phosphate-bearing ore. Inthe processrof Vthe present invention leached zone material, which is aclay-like material found in a layer between .the economic phosphatematrix and the silty-sand overburden in the Florida pebble phosphateelds, consisting predominantly of wavellite or pseudowavellite, togetherwith quartz, sand, kaolinite and uorapatite, is transported from themine to the processing plant either by dry or wet transportationsystems. If transported dry, the leached zone material is further dried,agglomerates disintegrated and then sized by mechanical or airclassication means to produce a fractionngf par.- ticles smaller thanbetween about `150 mesh and about 220 mesh standard screen size,depending upon theclassiication procedure in the ore dressing sectionvof-fthe plant. If the mined material is moved hydraulically Vthematerial is subjected to wet classification andthe undersized fractionsubjected to a thickening and filtering operation to produce a highsolids content slurry Vofthe order of about 30% to about 65% solidsbyweight. The solids of this slurry are preferablyY given a preliminarydrying step to reduce the moisture content thereof.

The dry solids whose small particle size fraction is the more valuableportion of the leached zone material because it contains roughly orbetter of the valuable minerals, will vary considerably in compositiondepending upon the area in which the ore is mined. Therefore, thefollowing description is given with reference vto a. leached zonematerial of which the following would be a representative assay obtainedby averaging about 200 samples.

Mineral Value Plus 200 Minus 200 Mesh Mesh 5. 44 14. 68 1. 68 26. 34 5.53 9. 19 l. 60 3. 11 UaOs. 0.0053 0.03 Acid Insoluble 83.08 38.V 56

For any one particular sample, the percentages of various minerals aresubject to considerable variation. For example, A1203 will vary fromabout 11% to about 35% in the minus 200 mesh fraction. With this varia;tion of one component there may be direct, inverse or randomvariation ofthe other components such as P205 and U303.

Leached zone material may be treated directly without any sizingoperation, with sulfuric acid, but in processing leached zone materialacidifying a minus 200 mesh fraction is preferred. j

n This normally dry minus 200 mesh fraction is ntimately mixed with.sulfuric acid in heavy duty mixing equipment such as a pug mill.Sulfuric acid is added to the mixer in the form of concentrated acid ofbetween about 70% and 98% sulfuric acid. This concentrated acid is addedin amounts giving a degree of aciduf lation which can be varieddepending upon the temperature level of heat treatment to followchemical and/or mineral composition of the feed and the length of timeof said heat treatment. Generally, sufficient acid is added to lgivebetween about 10% and about 90% acidulation, the exact range dependingupon the subsequent processing operations in handling of the leachliquors.rV If it is desired to have high uranium dissolution'andrelatively low phosphorousy and aluminum dissolution the preferredpercent acidulation` would befbetween about 20% raeefasoe 1 A and about50% while if relatively high recoveries of all components was desiredthe preferred percent acidulation would be between about 60% and about90%. Table I which follows illustrates the effect of percentVacidulation on a minus 200 mesh leached zone feed analyzing 30.4%A1203, 16.8% P205 and 0.026% U3O8. This ore required 1.04 poundssulfuric acid per pound feed for 100% acidulation. This ore was mixedwith sulfuric acid, heated three hours at 300 C., cooled and leachedwith water.

TABLE I Percent Solubilization for Percent Acidulation and ComponentIndicated Percent Acidulation Uranium Phosphorus Aluminum Variation inrecovery of mineral values from different source materials wasdetermined by mixing the required amount of sulfuric acid to give 40%acidulation with `each of eleven ore samples, heat treating the mix forabout three hours at 300 C., and leaching the cooled, heat treated mixeswith water. Results are shown in U3O8, P205 and A1203 recoveries insolution average 75%, 25% and 36%, respectively.

Percent acidulation referred to in this description is calculated on thebasis of the reaction of sulfuric acid rwith all of the iron, aluminum,calcium, magnesium and sodium or other signicant cationic constituentspresent in the leached zone material. In other words, 100% acidulationwould be the addition of that amount of sulfuric acid required tocompletely react with these components.

The elfect of increased temperatures, i.e., heat treating the ore attemperatures in the range of about 350 C. to about 725 C., for thosedegrees of acidulation which give relatively high recovery of uraniumand low recovery of phosphate and alumina is to improve the recovery ofuranium and lower the recoveries of phosphate and alumina. Temperaturescannot be raised indiscriminately, however, because above 600 C. someAsulfur dioxide is freed showing `evidence of decomposition of sulfatesand above 750 C., decomposition is so advanced that little solubilizingof any constituent results from the ore treatment.

At 40% acidulation of varioussamples of leached zone material, uraniumsolubilizing shows a relatively constant high recovery between about 250C. and about 700 C. 'with a slightly pronounced optimum generallyappearing at temperatures in the range between about `450 C.

and about 500 C. Over this temperature range of 250 C. to about 700 C.,bothphosphate and alumina recovery generally show a reduction withincrease in temperature with the net result of higher recovery ofuranium per unit of phosphorus and aluminum recovered. Composition ofores have a marked eiect upon response to acidulation and heat treatmentand it is to be understood that the above optimum range may not includethe optimum point for each and every sample of various ores as, forexample, of leached zone material or kaolinite.

Temperature of heat treatment and percent acidulation can beinterrelated to accomplish shifts in the relative amounts of uranium,aluminum and phosphate recovered. At any heat treatment temperature inexcess of 300 C., increase in percent acidulation to higher than about50% lowers the recovery of uranium. The same effect of lower uraniumrecovery is attained by elevation of the heat treatment temperatureabove 300 C., while maintaining a constant percent acidulation. Theabove conditions of increasing temperature While maintaining anyspecific percent acidulation also decreases the recovery of A1203 andP205. On the other hand, for any specific heat treatment temperature, anincrease in percent acidulation above 50% increased the recovery ofA1203 and P205.

Time of heat treatment is generally for a period of between aboutone-fourth hour and above six hours and preferably treated within aboutthree hours. This period of heat treatment may be varied considerablydepending upon the other variables discussed above, namely, percentacidulation and temperature of heat treatment. Normally, however, asmall increase in recovery is observed by increasing the heatingtreatment from about one to about three hours.

Slurry :formed in the pug mill may be heat treated in suitablecalcination equipment such as a hearth furnace or a rotary kiln wherethe solids move either co-current or countercurrent to the flow ofdrying gases. Drying may be accomplished by direct contact withcombustion gases or by indirect methods such as drying with hot airheated by heat exchange with hot combustion gases or by directapplication of heat to, for example, the rotary kiln.

Heat treated solids may be used directly in the process or stored untilneeded. Dry solids are crushed to produce a leaching operation feedhaving particles generally of less than about 8 mesh size. The groundsolids are then mixed with water or a weak water extract from previousprocessing, and slurried therewith in order to take up in aqueoussolution all of the water soluble values contained in the acidiiied heattreated leached zone material. This leaching operation is usuallycarried out in a multi-stage decantation or mixer filtering operation oftwo to four stages with the solids and water preferably movingcountercurrently through the stages.

When violent agitation is used during the leaching operation, heavystriation of the roasted frit takes place and iiltration rates aredecreased by a factor of -about 10 to 100.

A similar effect is obtained if the frit is finely ground beforeleaching. In neither case is the loss in filtration rate compensated byan increase in recovery of value. Mild agitation, therefore, isessential.

With mild mechanical treatment in the leaching step ltration andsettling rates are adversely affected as roasting temperature islowered. The character of the final frit varies lmarkedly withtemperature. At lower temperatures in the range of C. to about 250 C.,the roasted mass shows a lack of cohesion and disperses in water torelease large .amounts of fine material. Filtration rates for materialheat treated in this low range are reduced by a factor of three or more.The change in the character of the roasted frit occurs at differenttempera.- ture levels depending upon the percent acidulation. In

general, however, at the acidulation level of between about 20% andabout 40%, the adverse eifect of temperature of treatment generallyappears in the temperature range between about 200 C. and about 250 C.

1 Leached solution is adjusted to a speciic gravity in the range ofabout 1.15 to about 1.35 with about V1.2 to about 1.3 preferred. This isaccomplished by properwater addition duringY the leaching operation.With a. final liquor specific gravity of 1.3 and a iiltering temperatureof 65 C., lter rates of about l0 to about 60 gallons of liquor per hourper square foot of filter area are obtained. This is comparable to aiilter rate of about l to about 6 gallons of liquor per hour per squarefoot of iilter area when the leached zone material is not calcined.

I Filtration is carried out at temperatures as high as practical from anoperational point of view. Generally, the temperature of the slurry ismaintained at between about 30 C. and about 95 C., normally 65 C. A

So1ids-free leached solution when Vrecovered following processing underconditions where comparatively high recoveries of all components andparticularly alumina is obtained, is generally treated with ammoniumYcompounds such as ammonium sulfate and/ory ammonium acid sulfate tocrystallize out ammonium alum. Ammonium alum crystallization will varyin completeness of aluminum removal or in purity of product dependingupon such variables as Al2O3 concentration, time and temperature ofcrystallization and ratio of P205, NH4, S04, and Al ions.

Ammonium sulfate or ammonium bisulfate may be added to the solutionobtained from leached zone for crystallization of ammonium alum.Normally ammonia and sulfur values are added in excess of that requiredfor stoichiometric formation of ammonium alum.

When adding ammonium sulfate to give between about 1:1 and about 4:1NH4rAl mole ratios, the water content of the liquor has been found to bevery important. Normally as the NH4:A1 ratio is increased, under a givenset of operating conditions at about 25 C., the recovery of ammoniumalum as crystal will increase. This difference in recovery is leastapparent at about 60% water content. As the water concentration isincreased to about 80%, i.e., 20% dissolved solids by weight, the alumrecovery normally decreases. For example, when crystallizing at 30 C.from liquor having a 60% water concentration and a P2O5:Al2O3 mole ratioof about 0.6 in the solution, variation in ratios of NH4 to Al, from 1:1to 4:1 and when using ammonium sulfate, has little effect and aluminarecovery is of the order of 70 to 74%. At 80% water concentration on theother hand, at an NH4zAl ratio of 1, the recovery of alumina is reducedfrom about 70% to about 40%, the recovery of alumina increasing with anincrease in ratio until at a ratio NH41A1 of 4:1 the alumina recovery isonly reduced from about 71% to about 65%. In general when ammoniumbisulfate is substituted for ammonium sulfate, under the samecrystallizing conditions, alum recoveries are higher at a given NH4/A1ratio. At any waterV concentration up to 80%, the alum recoveries whenusing NH4/Al ratios between about 2 and about 4 show little or notendency toward reduction in alum recovery. When using ammonia bisulfateat 80% water concentration, and an NH4/Al ratio of 4:1 under identicalprecipitation conditions as in the case of ammonium sulfate previouslydiscussed, alumina recovery isof the order of 82%. When the sulfate ionin the precipitating liquor is increased above the NHiHSO.,= additionstage, i.e., addition of equal molar quantities of NH4HSO4 and H2804such that there is an excess of free sulfuric acid, alum recoveriesdecrease when NH4/Al ratios are increased, particularly as the watercontent of the precipitating liquor is decreased from about y80% andabout 60%. Under these conditions a NH4/Al ratio of about 1:1 sometimesissatisfactory to giveralum recov-V eries above 8.0%. I, n

I Under crystallizing conditions identical with those described above,as the P2O5/Al203 mole Vratio increases there is a decrease in alumrecovery. Solutions recovered from leached zone reaction with sulfuricacid give ratios between about 0.05 and about 1.0 normally between about0.5 and about 0.7 P2O5/Al202 mole ratios. Generally a crystallizationtemperature between about C. and about 35 C., preferably about 25 C., isused. crystallization time after addition of the desired compound, i.e.,ammonium bisulfate crystal, of less than six hours is used. Althoughlonger times will give somewhat higher yields, the additional equipmentrequired in general is not economically justified. Using combinations ofthe 'above variables alum yields between about 70% and about 97%,

, normally between about 80% and 85%, can be attained.

Upon removal ofthe crude ammonium alum from the mother liquor byseparation means such as filtration or centrifugingythe mother liquorthen is processed for subsequent recoveries of uranium and phosphorusvalues. Since the ammonium alum may be processed for recovery ofalumina, additional purification steps may be employed.

Suitability of alumina for use in the metallurgical aluminum isdetermined by P205, Fe203, SiO2, and other impurities. It has been foundthat these impurities will be removed by recrystallization of ammoniumalum.

Normally, two recrystallizations are required in order to remove thedesired amounts of these impurities. However, one recrystallization hasproven satisfactory in some cases. These recrystallizations normally areconducted in a continuous Oslo-Krystal ycrystallizer operating at about35 C. where alum crystals of between about minus 14 mesh and about plus60 mesh are produced.

If crude alum is recrystallized twice to reduce the P205, Fe202 andSiO-2 contentof the alum, the reduction in quantity of impurities isshown from the following table.

Ammonium alum recovered from the recrystallizing operations is reactedwith aqueous ammonia under conditions of very mild agitation and roomtemperature such that an alumina (Al2O3.XH2O) is produced.

Aqueous ammonia is mixed with the crystalline ammonium alum generallysuch that about %Y excess, of that required to neutralize the sulfatecomponent of aluminum sulfate, is added. Aqueous ammonia concentrationsbetween about 10% and about 30%, preferably about 20%, is used. Uponaddition of the reactants, the reaction slurry is generally agitated,i.e., by use of a slow revolving bailied rotary tube, for a period of upto about four hours. Substantially complete reaction occurs at 30minutes.` The resulting slurry is iltered and leached with about poundsof water per 100 pounds of ammonium alum initially used. Under optimumreaction conditions, fast iiltrations result. This pseudomorphous solidproduct, containing someammonia and sulfate impurities, `is heated tobetween about 900 C and about ll50 C. to recover alumina. Alternatively,if metallurgical grade alumina is not desired, the crude ammonium alummay be thermally decomposed at about 900 C. to produce an aluminadirectly.

Partially purified liquor normally free of suspende solids is subjectedto uranium recovery by use of solvent extraction as the preferredmodification. This solution is preferably iirst subjected to a reductionreactionA accomplished by electrolytic means or by chemical reactionwherein the solution is treated with metallic iron, etc., or other.reducingagents If the reducing agent is a solid, its removal isaccomplished through the use of a filter, centrifuge, cyclone or othersuitable separation devices'.- While'reduction is preferreisolvent.extraction f maybe employed uponsolution in an unreduced :as well vAmesh .standardy screen.

.ing the ollowingchemcal analysis: f l v .f

Percent Weight' as a' partially reduced state. After'removal of reducingCompnent agents,- they reducedl aqueous solution is; intimatelyA conf lA1203 t 30A taeted .or otherwise agitated with the lorganic' Solvent 5P205 1 16.8 extractant, Ugg@8 0.026 f Thisv extrlan'k is' made im? :oftw components the Ca() a5 i extractantand the vehicle or extender.l Theextractant d FeQ; 7 f 1,5 may be one or more of theortho and/or the pyrophos- "nf-nf "7"" n-n-nn phoric acid esters ofthe alkylmonohydricalcohols. vBoth l()v Afld msufbles (8102 m'm 7"- 30? i :the mono anddiesters,as Well as mixtures of the two are "Was mixed Wifh 96%.sulfuric? acl@ at a Tat? Off about 5.3

'uSefuL yhosphoc acid esters of octy or higher m0. pounds of acid'perlliOOJpounds oidry: sohds'and about f lecular weight alcohols'arepreferredsincezthey are ess .209 PQUDdS 0f Wat added- Pel' 10UPOundS Ofdfy 501]?15' water sembla The extender may be any oneormorbf:This amount of acid .constituted about 49% iacidulatlon. the commonorganic solveutsfsuchl as kerosene, benzene, l5 The 50h55' and liquidsWere' mlxedl-D'n'tflt10II mlXfz mineral'spirits, lcarbon tetrachloride.and the like. The ma@ about 80 fC- for# 18 hurs to Obtain dlssolutlomofconcentration of extractant in the extender: may: :QQHIPQBSDS- f f v t yWidely, for example, between about 5% andlabout 95%, The Slfl'l affe",CGOIIUE" *Was* @ached Wlh Wai?? 'at preferably between about and labout10%. yThe ,65 fr 30 mmutes, and 'filtered to :I 'ever slublhzgd "volumerasesofaqaeoussolutionteierganieseiventmay12e:rconstlfusntsandthmsplalbl .Cake dlsafded A lter f f vary withia'wide 1min, for' example,between about V1:1 rate @tf approxlmately mi lgallons lO f slurry-per.Square r f l and Aabout A48:11', preferablyI between about 3:1r andfoot 0f lter area Pel.' hour was attamedf Th@ resultlami about ;1iextract vat approximately v1.3 specific `gravity showed the The organicphase is treated ywith'aqueous hydrouoric following pdu'nds ofconstituents' :recovered perl A1000; f acid ori any othermixture-capableof precipitating and/ or Pounds Off mmus 20@ mesh ieachdZion@ feedf f v removing uraniumag for example, initheform :ofUF.: P205:plaga-ieg ..c. .M 108 The aqueous phase afterfextraction treatment isproc- A1203 .17.0. v essed lto recover the predominantly phosphorus,nitrogen U3O8 2.--.. Y- 0.18 l and'suliur values which? aff? in SQUOHHaHY ,Ilumbl' 'Of The: mineral digest solutionhavin-g a volume of aboutcompounds such as. ammonium sulfate, ammonium acid 30 239 gallonsWassubjectedm Contact withaboutlapguuds v v Sillfef metal 'phspha P205-Vhles iSUCh 3S PIQSFIOYC of powdered metallic iron and agitated forabouty 30 acid orpvmbirled :with variousimetal '@rsatoni val-ues lminutes, after. whichl :the solids were altered from .the f l l 1 asmentioned above. f l l 1 liquid; This liquid wasv then thoroughlycontacted with l i f v i MCher liquor: Sfpa'fed fl'Om G'll ammnlum alum'about 23 gallons of :an organic solvent composed of about l andPIOCeSSd 'fof Uamum' TeCOfVeIY 'then' iS :neutralized :35' I9Iparts .byvolume of kerosene and l part by volume of with ammonia forprecipitation of insoluble Phosphat@l y: a mixture er mondana:diesters:of-orthophesphorieacidf y y y I l l Compounds ContainingA1203, .F6203 and smania values f of seeetyi alcohol.' The' Contact wasmaintained foi-'1v1 i at: a PH 5110"@ :abol'tzp' 32,' Pfefeilly l'PH:21b0l14s5r l l about 2 minutes: inl each voizthe foursuccessive'counterj and ina continuous system. Upon separation of thein-I current s'tages.- The organic solventl was separated.fromy v v v ll l 50111512 phosphates Whr fasi fltfall rateS are Obtained 40 Contactwith the freshv aqueous phase and processed forA f f'theneutralizationly at'pH 4.5 has lbeendorre: inacontin@ l Tgeovei-yofaboutfog] pm'mds uranium fluoride gale@ (dry nous autogenous system, theresulting mother liquor conbasis), Containing 0.17 pounds Usos, byContact `with tains predominantly ammonium sulfate and monoammo- 15% HFsolution. Fresh organic was contacted with nium phosphate, exact ratio,concentration and quantities the aqueous phase, after three contactswith partially 4.0f each depending P011 the Prior PIOCCSSing employed.loaded (H308) organic, and upon separation of the phases This solutionis evaporated at boiling temperatures to a the aqueous phase wasprocessed to recover values conpoint where a rich fraction ofpredominantly ammonium tained in the aqueous phase. sulfate exists insolid phase. The resulting solution was adjusted to pH 7 with Uponseparation, near boiling temperatures, of the ammonium hydroxide (29%)and evaporated to dryness ammonium sulfate from the liquor phase, theresulting to recover a material suitable for agricultural conliquor iscooled to about 25 C. and monoammonium sumption. phosphate, containingsmall amounts of ammonium sul- Example II fate impurities, removed fromthe liquor phase. This liquor phase then may be cycled for subsequentrecovery 10%yplwagnntegs siszgeclhldcog i rcentaltng Vlus I/IonoammomumPhOFPhlte may 55 mesh standard screen. This leached zone mateia-l,havrys a 1re using .t e same type of c1rcu1t and phase mg the followingchemical analysis: system 1f a hlgher purity product 1s desired. C

Products such as ammonium aluminum sulfate, alumi- Omnnt' Percent byWelght num phosphate, iron phosphate and the like may be proc- P (2) 3304 essed in a number of diierent Ways, either to recover the U2 O5 16'sproducts in a purified form or to convert the products to Cos 0'026other forms such as aluminum, iron oxide and the like. F6203Alternatively, the aqueous phase with or Without alum 7n-".- u removaland uranium extraction may be treated with am- Ald mslubles (S102, db.)u 30'3 monia and dried directly to produce a fertilizer mate- 55 'wasmixed with 96% Summe acid at a rat? of about 53 rial containing ammoniumsulfate, ammonium phosphates ggunds f ald per 10g pounds of dry Sohdsand about and metal phosphates. The gure illustrates the above W tpoun-So Water ad -ed per 100 pounds 9fdry Solid-S described roce b a enbemgadded to a1d more complete mixing of solid Th p Ss means of a f lowdlagramand l1qu1d materials. This amount of acid constituted 1 einvenulon. will be further 1llustrated by the fol- 70 about 49%acl-mation t @Wmg eKamp es- The solids and liquids were mixed in a pugmill run- Example 1 lfllllg '131 about 11p-Ifn- FDlie heavy pastedischarged rom e u m1 was roaste t Dry leached zone material vas airsized to produce a proximatel/ 00" C. for three housa temperature of ap.1000 pound fraction of material passing Athrough a 200 The roastedmaterial after cooling was leachedwith This leached .zone material, hav

asoaso water at 65 C. for 30 minutes and filtered to recover solubilizedconstituents and the insoluble cake discarded. A lter rate ofapproximately 40 gallons of slurry per square foot of lter area per hourwas attained. The resultant extract at approximately 1.3 specic gravityshowed the following pounds of constituents recovered per 1000 pounds ofminus 200 mesh leached zone feed.

The mineral digest solution having a volume of about 200 gallons wassubjected to contact with about pounds of powdered metallic iron andagitated for about 30 min- Y utes, after which the solids were filteredfrom the liquid. This liquid was then thoroughly contacted with about 20gallons of an organic solvent composed of about nine parts by volume ofkerosene and one part by volume of a mixture of mono and diesters ofortho phosphoric acid of iso octyl alcohol. The contact was maintainedfor about two minutes in each of the four successive countercurrentstages. The organic solvent was separated from contact with the freshaqueous phase and processed for recovery of about 0.38 pound uraniumiluoride cake (dry basis), containing 0.23 pound USOS, by contact withHF solution. Fresh organic was contacted with the aqueous phase, afterthree prior contacts with partially loaded (U3O8) organic, and uponseparation of the phases the aqueous was then processed to recovervalues contained in the aqueous phase.

The resulting solution was adjusted to pH 7 with arnmonium hydroxide(29%) and evaporated to dryness to recover a material suitable foragricultural consumption.

4 Example III Dry leached zone material was air sized to produce a 1000pound fraction of material passing through a 200 mesh standard screen.This leached zone material, hav-V ing the following chemical analysis:

was mixed with 96% sulfuric acid at a rate of about 53 pounds of acidper 100 pounds of dry solids and about pounds of Water added per 100pounds of dry solids, Water being added to aid moreV complete mixing ofsolid and liquid materials. This amount of acid constituted about 49%acidulation.

The solids and liquids were mixed in a pug mill running at about 120r.p.m. The heavy paste discharged lfrom the pug mill was roasted at atemperature of approximately 500 C. for one hour. f

The roastedmaterial after Ycoolingy was leached with water at 65 C.for'30 minutesv and filtered to recover solubilized constituents and theinsoluble cake discarded. A filter rate of -approximatelyQSOgallonsno'fslurry per square foot of lilter area per hour was attained.The resultant extract at 'approximately 1.3 specific gravity showed thefollowing-pounds of constituents recovered per 1000 pounds of minus,20,0 mesh leached zone feed:

P205 62 A1203 146 Usos 0.23

The mineral digest solution having'a volume of about 150 gallons wassubjected to contact with about 10 poundsY of powdered metallic iron andagitated for about 30 minutes, after which the solids were filtered fromthe liquid.y Ilhis liquid was then thoroughly contacted with about 15gallons of van organic solvent composed of about nine parts by'volume ofkerosene and one part by volume of f of iso octyl alcohol. The contactwas maintained for about two minutes in each of the four successivecountercurrent stages. The organic solvent was separated from contactwith the fresh aqueous phase and processed for recovery of about 0.36pound uranium uoride cake (dry basis), containing 0.22 pound U3O2, bycontact with 15% HF solution. Fresh organic `was contacted with theaqueous phase, after three prior contacts with partially loaded (U3O3)organic, and upon separation of the phases the aqueous phase wasprocessed to recover values contained in the aqueous phase.

The resulting'solution was adjusted to pH 7 with am# monium hydroxide(29%) and evaporated to dryness to recover a material suitable foragricultural consumption.

. Example IV Dry leached zone materialwas air sized Yto produce a 1000pound fraction of material passing through a 200 mesh standard screen.This leached zone material, having the following chemical analysis:

was mixed with 96% sulfuric acid at a rate of about 81 pounds of acidper 100 pounds of dry solids and about 20 pounds of water added per 100pounds of dry solids. This amount of acid constituted about 75%acidulation.

The solids and liquids were mixed in a pug mill running at about r.p.m.The heavy paste discharged from the pug mill was roasted at atemperature of approximately 300 C. for three hours.

The roasted material after cooling was leached with water at 65 C. for30 minutes and filtered to recover solubilized constituents and theinsoluble cake discarded. A lter rate of approximately 35 gallons ofslurry per square foot of filter area per hour was attained. Theresultant extract at approximately 1.3 specific gravity showed thefollowing pounds of constituents recovered per 1000 pounds of minus 200mesh leached zone feed.

P205 A120a 250 U308 0.230

The solids removed in the ltering operation and washing were dried at110 C. to give 530 pounds of dry cake having the following composition:

Constituent: Percent by weight P205 6.6 A1203 10.1 U3O2 L 0.006 S04 4.6

Constituent: Percent by weight (dry basis) I A1203 9.98 P205 0.181 Fe2020.013 SiO2 0.013 S04 39.7 NH4"' 4.33

were recovered. crude ammonium alum crystal was Twice RecrystallizedAlum Recrystal- Impurty as Parts per Million lized Alum Component:Percent by weight A1203 12.8 S04 19.5 NH2 7.28 H20 (110 C.) 55.3

were recovered. These solids were fired for one hour at 1150 C. to give11.8 pounds of aluminum of sufcient purity and water sorption propertiesthat it could be used in the electrolytic production of high gradealuminum metal.

Aqueous solution recovered from the separation of the crude ammoniumalum was adjusted to a sulfate composition of 220 grams per liter togive about 206 gallons of solution having a specific gravity of 1.30.This solution was subjected to contact with about pounds of powderedmetallic iron and agitated for about 30 minutes, after which the solidswere filtered from the liquid. This liquid was then thoroughly contactedwith about 20 gallons of an organic solvent composed of about nine partsby volume of kerosene and one part by volume of a mixture of mono anddiesters of ortho phosphoric acid of iso octyl alcohol. The contact wasmaintained for about two minutes in each of the four successivecountercurrent stages. The organic solvent was separated from contactwith the fresh aqueous phase and processed for recovery of about 0.36pound uranium fluoride cake (dry basis), containing 0.22 pound U3O2, bycontact with 15% HF solution. Fresh organic was contacted with theaqueous phase, after three prior contacts with partially loaded (U308)organic, and upon separation of the phases the aqueous was thenprocessed to recover values contained in the aqueous phase.

The resulting solution heated to 100 F. was adjusted to pH 4.5 in acontinuous system with anhydrous ammonia gas, held at this pH for about30 minutes, filtered on a drum filter to give a filtering rate of 26gallons per hour per square foot of filtrate, the wet cake washed withan equal volume of water and this cake dried to give a product, suitablefor use in agricultural applications, of the following chemicalcomposition:

Component: Percent by weight P205 (total) 43.8 P205 (citrate soluble)42.85 P205 (water soluble) 4.15 A1203 18.35 NH2 7.22 Fe2O3 4.77 F 3.80

CaO 0.74 S04 1.28

12J whichon a dry `basis gave 207 pounds 4of product. Solutions from theltering and washing of the aluminum phosphate fertilizer material werecombined and evaporated at atmospheric pressure to give a liquor phasecomposition of,

50% NH4H2PO4 i 25% (NH4)2S04 then ltered hot, water added to thesolution and cooled to 25 C. to crystallize NH4H2PO4 and leave a liquorphase composition, in equilibrium with the solid phase at 25 C., of

37% (NH4)2S04 which was present at a weight ratio of crystallized solidsto liquor ratio of 30 to 80. Upon filtering at 25 Cfthe liquor phase wascycled for mixing with fresh liquor, after removal of the aluminumphosphate fertilizer material, and subsequent further recoveries of thevalues contained in this portion of the liquor. Approximately 680 poundsof material largely as ammonium sulfate and pounds of material largelyas monoammonium phosphate were recovered upon attainment of equilibriumon cycling of the liquor which was recovered upon separation ofmonoammonium phosphate.

As was illustrated in Tables I and II comparatively high uraniumrecovery and low phosphorus and aluminum recoveries can be obtained byproper selection of percent acidulation when using heat treatingtemperatures in the range of about 250 C. to about 350 C. Even higheruranium recoveries per unit of phosphorus and aluminum are possible byincreasing the heat treating temperature from about 350 C. to about 700C. while maintaining 40% acidulation. This has been illustrated in TableIII following. In these cases sulfuric acid was added to dry minus 200mesh leached zone, heated one hour at the temperature indicated, cooledand extracted 30 minutes at 95 C. Percent of original com,- ponentssolubilized are shown for feeds K, W and M. Feed W was determinedlto belargely wavellite and silica sand, feed M was a mixed ore ofpseudowavellite wavellite, kaolinite and silica sand, and feed K waslargely kaolin and silica sand. AA sample containing largelypseudowavellite and silica sand gave percent solubilization intermediatebetween feeds M and W and have been omitted.

TABLE III Percent solubilized Component Extracted at Tem erature f o o.Indicated p rom Feed Indicated IKK" NMI' HWI,

(1) Usos 55 70 82 16 66 62 (OU 48 40 a s 83 a ia; i i3 3 a 93 (2) PlO-ig (3) A1203 38 28 40 (D) 700 C.

(1) UBOB 83 46 77 a a .s 2 8 2 a 3 (E) 800:10. 6 10 20 (2) }Less than 5%of au components.

The decrease of values above about 700 C. is attrib.- uted to the lossof sulfur in the `heating operation.

Having thus described our invention what we claim is:

1. In a method of recovering mineral values from low grade phosphateore, the steps which comprise acidulating the ore Ywith'sulfurica`cid`under conditions of strong agitation'to'forma pastyslurry, heattreatingvtheV acid mix at temperatures inthe range between about 250 C.and about 700 C. while drying the acid mix to produce dry solidstherefrom, and leaching the water soluble reaction products from saiddry solids, whereby a leach solution is obtained containing phosphorus,aluminum and uranium values. a

2. In a method of recovering mineral y values from leached zonematerial, the steps which comprise classifying the ore to recover aminus 200 mesh fraction, acidulating the leached zone fraction withsulfuric acid under conditions of strong agitation to form a pastyslurry, heat treating the acid mix at temperatures in the range betweenabout 250 C. and about 700 C. while drying the acid mixto produce drysolids therefrom, and leaching the water soluble reaction products fromsaid dry solids, whereby a Yleach solution is obtained containingphosphorus, aluminum and uranium values. n

3. A method of recovering mineral values from leached zone materialwhich comprises classifying the ore to recover a minus 200 meshfraction, acidulating the fraction with sulfuric acid under conditionsof strong agitation to lform a pasty slurry, said sulfuric acid beingadded in an amount to produce between about and about 90% acidulation,heat treating the acid mix at temperatures in the range between about250 C. and about 700 C. While drying the acid mix to produce dry solidstherefrom, leaching the vwater soluble reaction products from said drysolids and recovering from the leached solution at leastrone constituentselected from the group of phosphorus, aluminum and uranium.

4. A method of recovering mineral values from leached zone materialwhich comprises classifying the ore to recover a minus 200 meshfraction, acidulating the fraction with sulfuric acid under conditionsof strong agitation to form a pasty slurry, said sulfuric acid beingadded in an amount to produce between about 60% and about 90%acidulation, heat treating the acid mix at temperatures in the rangebetween about 250 C. and about 700 C. while drying the acid mix toproduce dry solids therefrom, leaching the water soluble reactionproducts from said dry solids and recovering from the leached solutionat least one constituent selected from the group of phosphorus, aluminumand uranium.

5. A method of recovering mineral values from leached zone materialwhich comprises classifying the ore to recover a minus 200 meshfraction, acidulating the leached zone fraction with sulfuric acid underconditions of strong agitation to form a pasty slurry, heat treating theacid mix at temperatures in the range between about 250 C. and about 700C. while drying the acid mix to produce dry solids therefrom, for aperiod between about one-half hour and about six hours, leaching thewater soluble reaction products from said dry solids and recovering fromthe leached solution at least one constituent selected from the group ofphosphorus, aluminum and uranium.

6. A method of recovering mineral values from leached zone materialwhich comprises classifying the ore to recover a minus 200 meshfraction, acidulating the leached zone fraction with sulfuric acid underconditions of strong agitation to form a pasty slurry, heat treating theacid mix at temperatures in the range between about 250 C. and about 700C. while drying the acid mix to produce dry solids therefrom, for aperiod between about one and about three hours, leaching the watersoluble reaction products from said dry solids and recovering from theleached solution at least one constituent selected from the group ofphosphorus, aluminum and uranium.

v7. A method of recovering mineral values from leached zone materialwhich comprises classifying the ore to recover a minus 200 meshfraction, acidulating the ore with sulfuric acid under conditions ofstrong agitation to form a pasty slurry, aging the acid mix for a periodof about fourteen to Vabout thirty days, heat treatingY Vthe acid attemperatures in the range between about 250 C. and Vabout, 700l C. whiledrying the acid mix to produce dry solids therefrom, leaching the watersoluble reaction products'from said Vdry solids and recoveringfrom theleached solution at least one constituent selected from the group ofphosphorus, aluminum and uranium.

i 8. A method of recovering mineral values from leached zone materialwhich comprises classifying the ore to recover a'minus '200 vmeshfraction, acidulating the leached zone fraction with sulfuric acid underconditions of strong agitation to form a pasty` slurry, heat treatingthe acid mix at temperatures in the range between about 250 C. and about700 C.`while drying the acid mix to produce drysolids therefrom,leaching the water soluble reaction products from said dry solids,separating the solution of reaction products from insoluble material,reacting the solution with Ysulfates of ammonia to precipitate alum,separating the ammonium alum precipitate and recovering from theresultant solution at least one constituent selected from the group ofphosphorus and uranium. Y Y

9. A method of recovering mineral values from leached zone materialwhich comprises classifying the ore to recover a minus 200 meshfraction, acidulating the leached zone fraction with sulfuric acid underconditions of strong agitation to form a pasty slurry, heat treating theacid mix at vtemperatures in the range between about 250 C. and about450 C. while drying the acid mix to produce dry solids therefrom,leaching the water soluble reaction products from said dry solids,separating leached Asolution from insoluble material, reacting theleached solution with sulfate of ammonia to precipitate ammonium alumeiecting at least a partial reduction of the uranium present in theresultant aqueous liquor to the quadrivalent stage, contacting thereduced solution with an organic solvent containing a phosphoric acidester of an alkyl monohydric alcohol, separating the organic phase fromthe aqueous phase, treating the organic phase with sucient aqueous HF toprecipitate uranium tetrafluoride and segregating the UF4 precipitate.

10. A method of recovering mineral values from leached zone materialwhich comprises classifying the ore to recover a minus 200 meshfraction, acidulating the leached zone fraction with sulfuric acid underconditions of strong agitation to form a pasty slurry, heat treating theacid mix at temperatures in the range between about 250 C. and about 450C. while drying the acid mix to produce dry solids therefrom, leachingthe water soluble reaction products from said dry solids, separatingleached solution from insoluble material, reacting the leached solutionwith sulfate of ammonia to precipitate ammonium alum, effecting atleasta partial reduction of the uranium present in the resultant aqueousliquor to the quadrivalent stage, contacting the reduced solution withan organic solvent containing a phosphoric acid ester of an alkylmonohydric alcohol, separating the organic phase from the aqueoussolution phase, treating the organic phase with sufficient aqueous HF toprecipitate uranium tetrafluoride, segregating the UF., precipitate andammoniating the recovered aqueous solution phase.

ll. A method of recovering mineral -values from leached zone materialwhich comprises classifying the ore to recover a minus 200-meshfraction, acidulating the leached zone fraction with sulfuric acid underconditions of strong agitation to form a pasty slurry, heat treating theacid mix at temperatures in the range between about 250 C. and about 700C. while drying the acid mix to produce dry solids therefrom, leachingthe water soluble reaction products from said dry solids, separatingleached solution from insoluble material, reacting the leached solutionwith sulfate of ammonia to precipitate ammonium alum, effecting at leasta partial reduction of the uranium present in the resultant aqueousliquor to the quadrivalent stage, contacting the reduced solutionv withan organic solvent containing a phosphoric racid ester ot an alkylInonohydric alcohol, separating `the organic phase from theaqueous'phase, treating the oryganic phase with sufcient aqueous HF to'precipitate uranium tetrafluoride, segregating the UFQ precipitate, ad-

ore'to recover a minus 200 mesh fraction, adding to the fleached rzonematerial which comprises classifying the leached zone fraction 96%sulfuric acid in a quantity of f about 53 pounds of acid per 10() poundsof dry solids,y

adding about 20 pounds of vwater to the acidied mix and agitating themix ina pug mill toy form afheavyy paste, heat treating the mixatatemperature of approxi-v mately 300? C. for' three hours vwhile`drying the mix to produce dry solids therefrom,'leachingy said ydrysolids with water to produce an .extract of approximately 1.3 specificgravity, contacting the extract with powdered metallic iron for about 30minutes to reduce uraniumto the tetravalent state, separating the ironsolids, contacting the reduced extract with a vkeroseneextender mixturevof mono and diesters of ortho phosphoric acid, separating the aqueoussolution and organic ester phases, mixing the organic ester phase with15% hydrofluoric acid to precipitate uranium tetrauoride, separating anddrying the UFQ, adjusting thepHof the aqueous phase to about f 7with'ammonium hydroxide and evaporating the neutralized material todryness.

13.- Ay -method .of recovering mineral values from rore to recover aminus 200 mesh fraction, adding tothe leached zone fraction 96%sulfuricy acid at a rate of rabout 53 pounds, of acid per 100, pounds ofdry solids, t adding about 20 pounds of Water to the mix and agitatingthe mix kin a pug mill to form a heavy paste, heattreatf ing they mixata temperature ofvapproximately 500 C.

for one hour while drying the mix to produce dry solids therefrom,leaching said dry solids with Water to produce yan extract ofapproximately 1.3 vspecific gravity, contacting the extract withpowdered metallic iron ,for about 30 minutes to reduce uranium to thetetravalent state, separating the iron solids, contacting the reducedextract -f with a kerosene extender mixture of mono and ydiesters ofortho phosphoric acid, separating the vaqueous solution andyorganic'ester phases, mixing the organic ester phase with 15%hydrofluoricacid to precipitate ,uraniumtetrwv uori'de, separating anddrying the UF4, adjusting the pH ofthe aqueous phase to about 7 withammonium hydrox f ide and evaporating the neutralized material toydryness. References -Cited n the le of'this patent UNITED STATES PATENTSn y 1,098,282. McCoy May 26, 1914 2,769,686 McCullough et al. Nom-6,1956y i McCulloughet al. .f lan. 7, 1958 OTHER REFERENCES RMO-2030, AECDocument, sentis, 1954.

...RMO-2,032, AEC DocumenQSept. 13, 1954, pp. 6,v

RMO-2041, AEC Document, Feb, 28, 1955,v PP. 6, 7..

1. IN A METHOD OF RECOVERING MINERAL VALUES FROM LOW GRADE PHOSPHATEORE, THE STEPS WHICH COMPRISE ACIDULATING THE ORE WITH SULFURIC ACIDUNDER CONDITIONS OF STRONG AGITATION TO FORM A PASTRY SLURRY, HEATTREATING THE ACID MIX AT TEMPERATURES IN THE RANGE BETWEEN ABOUT 250*C.AND ABOUT 700*C. WHILE DRYING THE ACID MIX TO PRODUCE DRY